Uranium recovery process



Feb. 10, 1959 R. H. BAlLEs ETAL 2,873,165

URANIUM RECOVERY PROCESS Filed May 26, 195o 3 sheets-sheet 1 Petz. 11o,1959 R. H. BAILES ETAL URANIUM RECOVERY PROCESS Filed May 26, 1950 PERGENT URANIUM RECOVERY IOO 3 Sheets-Sheet 2 A O L53 EQU|VALENTS F/l. AL75 EQUIVALEN-l- S F/l.

CATION EQUIVALENTS FL`UORlDE EQUIVALENTS NVENTORS. RICHARD H. BA/LESROBERT S. OLSON HERBERT 0. KERL/NGE? By RAY 5. LONG TTORNEK feb. 10,1959 R. H. BAILEs rs1-A1.

URANIUM RECOVERY PROCESS Filed may 2e, 195o R E .3 BNG OMUN SIG .tTALLN.. NBOHO .n \l, E .EL S .m WHSK. Q- m 4 .t RRTY e O AE A e MRM E 3m H Y m B O nlv O E 8 m T u o le N H250., V H2504 E D. C A

A T TORNEY.

F, IC

.- v Fei. 10, 1959 r as nal products evsiiitab-le for processing by themethod` of the invention. l i `illus.traltively, such raffscilution-kmay have the composi- `URANIUM RECOVERY PRocEss tion indicated in 'the.tellbwislftsbler Richard LNI. Bailes, Walnut Creek, Ray S.Longfvalleio, Egberts. Olsdm San ErancscoQnd HerbrtOfKerf linger,Berkeley, Calif., as'signorsx, `by"mesne assignments, toV `the .United-`States of America as represented by the United StatesAtomic EnergyCommissioni Application May 26, 1950, Serial No. 164,566

14 crains. (c1. zel-14.5)

d This invention relates to processes for recovering uranium fromsolutions and, `more particularly, `relates to the recovery of uraniumfrom phosphoric' 'acidlor other phosphatic solutions.

`The presence oflow concentrations of uranium in certain`phsphaticoreshas beenknown forlsome time but fthe direct recovery Aofuranium from this source has llQttben QC.Onlvlllsally` feasible dus t0the 10W concentrationpresent, SinceA these ores are pres-V onlyv P Hlledinthe Preparation 0f Superphosphate and phpsphoric acid-l in `processes`wherein sulfuric acid` is etuplgyedtonuleaching the phosphatic materials`from the ore herginthe uraniumvis leached intothe recfphosphaticnsolutions, there existsdfa posgctluvering. this low` grade`uram'lum from` the of uranium from th`e"phosphatic Vsolution complicatedbeth `bythe 19u concentration thereof and the high Cenu ccntration'y 'ofphosphate in thel solution in `which'solution .the uranium is highlysoluble due.- to complexing phenomena." Y l owfit lbeen iound thatylowfconcentrationsof .isomer bsfesovslectfrem crude phosphoric acid er?other ias'fific phqsphatc Sfluies by reducing `.the uranium to thetetrai/alent state and precipitatingi .the as` awtdiuoride and, further,that 4the uranium' methods 1.1.1.0..` fully' described" hereinafter..

,Ascetdiesly it' is 211,1 Object ef the invention to provide a methodfor recovering uranium from phosphoric acid or` other phosphaticsolutions.

AAnother object of the invention is to provide a fluoride CII However,solutions diteriug very considerably from that indicated i A relativedcnc'ent-rationsdorin the types of itms pre,Y y als be satisfactorilyprocessed.

With refer ce t thetou hee't o f `Fig. Land in accordance 'inventionyalpl'iosphafte or phosphoric acidV solution offfthe `cljra'rafct,erdescribed is subjected to reduction, whereby all'of' the uraniumcontrzrtinedl therein vis reduced to the tetravalent state. Thereduction of the uranium to the tetravalent state may be accomplishedutilizing metallic reducing agents such asiron, zinc, zinc amalgam,`aluminum or lead, or byV using reducing agents such as hydrazine salts,hypophosphorous acid or hydroxylamine hydrochloride or'by electrolyticmethods. The Yrequirements for the reducing'agent are no t critical anda majorityof'the'reducing agents having a suiciently high reducingpotential appear to besuitable ,for the purpose. l.'l`-he..pro"gress` ofthe reduction' is easily followed by observing; the. electromotivepotential developed between platinum and saturated calomel electrodeslimmersed in the solution. Reduction is substantially connplif@v .when.the potential is lowered to about +O.1,.yolt.

l `Policy/ving reduction, the solution is treated to coprecipitate;thereduced Auranium as a fluoride together with other iIlSQlllble iuoridesthereby achieving a subi stantially `complete removal of even very smallquantipreipitation 'rne'thod for recovering uranium froripliosphoricacid or'other'phosphatic "solutionsl l "A furtherobject of'the'iventionis to provide uranium `ieovery processes vwherein a fluoridepreciisita'te ontaining' uranium is deuor'inated andthe uranium isA-recovered fromthe defluorinated precipitate.

stillturther 4object of. the invention is to provide uranium .recoveryprocesses wherein uranium is 'leached from. a.uoridefprecipitatefand theuranium is recovered from the `.leach solution. Y

kOther robjects and advantages of the invention will become.apparentffromlthe following description taken in conjunction -with theaccompanying drawing `of which: Eigurefl is' a how sheet of theprocesses of the invention; i d d' fFig. 2is a graphical illustration ofthe ndependence of uraniumV recovery upon precipitant ratios andamounts; and

Fig. 3 is aigraphical illustration of the dependence of the leachingcharacteristics of deliuorinated ndcalcined uoride vprecipitates .upontime and temperaturefof icalcination. Relatively concentratedphosphoricacid or phosphatc solutions containing uranium in" lowconcentrations fs'uch as lare obtained in"v`arious.phosphatefertilizerand metallurgial y, ,progesses g. as byproducts; intermediate' :productsties of. uranium from the phosphatic solution. event :dthiat theI ltonis `d ei'itent in lcoprecipzitant materials, fa sulic .q offasolubley compound of the coprecipitant `isy supp Y to.fthe'solution andthen a SolublF-fluoideger hydefl c acid .is added t0 .coprecipitats thereduced. ,grausam with. the 1Cor-recipient msterial. Calcium carbonate,calcium oxide, hydrated calciutn o aide, andbariuml carbonate have beenfound In the satisfactbryfftoE supply the required coprecipitant matefrials; however, other `materials@may also be used as will` becomeapparentffrom theff'llowingdiscussion. vVarious solubleuoriddes, suchasf'sodiumhuoride, potassium fluoride, ,bariurntluoride and ammoniumfluoride vand other fluorideswhich `fdurnislr'al Vconsiderable fluorideion concentration tohefsolution, may be employed in the precipitation,also hydrouoric acid (hydrogen uoride) is completely satisfactory. "['heratio of cation equivalents ,to ttuoride equivalents present in the acidand the total amount'of precipitants present determine to a critical`degree the amount of uranium which is recovered in the precipitate.`Fig. `2` graphically illustrates the close and critical relationshipwhich these factors bear to the amount offuranium recoveredL` Themechanismof .the precipitation of uranium from d thephosphnjic,`acidrsolution is not completely understood. AIt has been noted that thepresence of aluminum in the solution ygreatly increases the recovery ofuranium. In the presence of; aluminum va cryolite type of compoundlprecipitatesfwhich compound is preferentially precipitated rather thanuranous i fluoride. The cryolite s compound which precipitates appearsto have the general formula X (AIFBL, wherein X may represent variousproportions and combinations of radicals such as uranous uranium, Mg,Ca, Na, K, NH4 and others. The existence of compounds such as `MgaUllF):and Ca3(AlF)2 in the precipitate has been indicated by X-ray diffractionstudies. Moreover, a; mixture of variousV ,otherfinsoluble tluorides maybe precipitated along with the insoluble uranium compound. Some of theiluoride originally present in the acid and thatadded during theprecipitation may be recovered Vfrom the precipitate -andfrom thesolution forn recyclingby methods more.fully..describedhereinafter.4

The uorideprecipitate Acontaining uranium is. separated fromthegsolution, bylltering and the uranium is recovered' from theprecipitate by one of several valternative methods and the ltrate is.treatedfor recovery of the uoride or/and is returned to .theplant fornormal PfC6SS1Ug- A typical precipitate derived from such a solutionusing calcium carbonate and'hydrogenvuoride as the precipitants may havea composition determined by analysis approximately as follows: However,it will be api preciated that the composition 'of such a precipitate mayvary considerably' from that indicated dependent upon the composition ofthe original acid 'and conditions -of precipitation. -v-

Percent U3O8 0.25 Phosphate 1.95 Sulfate v 0.85 Ca. A1 v e A 8.6 Mg 4.95F 47.0

Emission spectrographic analysis discloses Y thatA other materials maybe present as follows:

Percent Fe 0.5-1.0 Na '0.5-,1-10 Si "0.05-0Ql0" V 0.050.10 Cu f01001-001 Cr l 0.001-001 Mn 0.00l-0.0l Ti 0.0014001 In general themethods of uranium recovery whichare employed may be classed in twogroups; one, in which the uranium is recovered from the precipitatefollowing defluorination of the precipitate; or, two, in which theuranium is recovered from the precipitate prior to defluorination. i

In the methods of the iirst classification, sulfuric acid is added tothe precipitate in stoichiometric. excess relative to the quantity offluoride present and hydrogen liuoride is vaporized from the precipitateand recovered in order that it may be recycled. `Alternatively, ammoniumsulfate is added to the precipitate and ammo nium fluoride issublimed'from the precipitate and recovered for recycling. Then thedelluorinated precipitate is calcined and leached with water or aVsolution such as dilute hydrochloric, nitric or sulfuric acid solutionsand the titanium is finally recovered from the leach solution.

The temperature to which the precipitate is heated in driving oftl theiiuoride atects the conditions which must be employed during theleaching. If the'precipitate, following treatment with sulfuric acid, isheated to a temperature below about 800 C., for 1% to 2 hours theprecipitate produces an acidic solution on contact with water sincealuminum sulfate appears to be formed by this treatment which aluminumsulfate produces a sufliciently acid solution to extract the uranium.However, if the heating time is prolonged or the temperature is.increased to about SOO-1000" C., alkaline products prey 4 dominate inthe heated precipitate and au acidic solution must be employed to leachthe uranium. Moreover, the solubility of various components of theprecipitate is markedlyv affected by the time and temperature of heating; Ilittle aluminum and/ or magnesium being leached from the moreintensively heated materials.

The composition of a leach liquor typical of those which are derived byleaching a iluoride precipitate, which was heated to a temperature of800 to 1000" C. for 1% to 3 hours with a slight excess vof sulfuricacid, employing 1% sulfuric acid is indicated below in Table 2., About80-90% of the uranium, none of the aluminum, and only one-third of themagnesium was'leached. l I

' calcination and leach reagent are graphically indicated in e this typefollows:

Fig. 3. The time-temperaturer curve (A) of'deuorination vare plottedwith respectv to the leach'liquorpH (reading upwards) curve (B);percentage-ofuran1um leached,'curve (C); percentage of aluminum'leached,curve (D); and percentage of Mg leached, curve (E).

Optimum conditions for a sulfuric acidleach are as follows: y0.2 N orabout 1% sulfuric acid as the leach reagentminus 40 mesh size of theprecipitate; and a liquidto solid ratio of about 2.5: 1.

The uranium may berecovered from the leach liquor by two methods:

(l) Precipitation of uranyl phosphate by pH adjustment, yielding 99.9%uranium recoveries ofl a material which after ignition containsuraniumequivalent to about 10% U3O8, employing amounts of sodiumhydroxide equivalent. to about 8 to 16 gramsv per gram of USOSprecipitated.

(2) By reduction of the leach liquor with sodium hydrosultite, or othersuitable reducing agents, 99+% of the uranium is precipitated in amaterial containing about 10 to 20% UaOg. An analysis of a typicalprecipitate of Percent 13.8 Phosphate 30.3 Sulfate 1 Ca n Y Mg In themethods of the second classication, hot sodium carbonate is employed toleach the uranium from the fluoride precipitate with great' etiiciencyWhile ammonium carbonate may also be employed if somewhat less leachingeiiciency can be tolerated. The leaching eiciency of sodium carbonatesolutions drops off very rapidly below a critical pH value between about8.5.and '9.0.

o A decrease in eiiiciency isnoted with increasing solid to liquidratios and a marked increase if the temperature at which the leaching isperformed is iucreasedjtonear the boilingA point. ACarbonate leaches alloftheV sulfate; about 0.1% of the phosphate, 0.8% of the aluminum and 2%of the fluoride from the precipitate in addition to the'd uranium. Y

. sodiumcarbonate leach solution preferably by neutralizing thesolution' with hydrochloric acid and rprecipitating the, uraniumy in japurified' form with carbonate-free S arnrnonia as ammonium vdiuranate.If there is present a considerable concentration of phosphate in thesolution, uranyl phosphate or a mixture of uranyl phosphate and ammoniumdiuranate is precipitated. Precipitated ammonium diuranate is thencalcined to yield high grade U30. The leached vprecipitate may then betreated to recover the uorine or is discarded.

Alternatively, in the second classication of methods, acidic leaching-ofthe fluoride precipitate may be employed without washing theprecipitate. Dilute aqueous solutions of hydrochloric, nitric, citricacid and less effectively tartaric, acetic and-boric acids are used toleach the uranium from the precipitate with the uranium concentrationbeing enhanced by successively leaching several portions of precipitatewith the same solution and uraniumv is recovered from the acid leachsolution by precipitation methods which further increase the purity ofthe material.

Acid leaching of the precipitate has been found to be influenced by thefollowing variables as exempliiied when dilute hydrochloric acid isemployed; concentration of acid, time of leaching, temperature, solid toliquid ratio and phosphate concentration in the uoride precipitate.Optimum conditions for leaching more than 95% of the uranium from a dryfluoride precipitate require `a 30 to 35 minute leach with agitation at95 C. with 5 to l0 ml. of 5% hydrochloric acid per gram of uorideprecipitate. Lowering the leaching temperature to 25 C. decreases theuranium recovery to 75% and a live minute leach yields nly"30% uraniumrecovery. Leach times beyond 30 minutesshow'no improvement in uraniumrecovery. When acid concentrations above about are employed the loss offluoride from the precipitate becomes large and lower concentrationsdecrease the uranium recovery. Markedly decreased uranium recovery isnoted with lower solid to liquid ratios while higher liquid to solidratios show only a'slight improvement.

`Reduction of the uranium inthe leach solution with variousreducingagents causesthe formation of a uranous kphosphate precipate with thephosphate which is leached into the hydrochloric `acid solution. Sodiumhydrosulfte, added to the solution as a reductant, precipitates amaterial containing uranium equivalent to about 55 to`65% U3O8 and about30% of phosphate.

Particular details of the processesA of the invention will becomeapparent from a consideration of the following illustrativeexamples ofthe operation of processes in accordance with the invention:

EXAMPLE IA 17:1 `liters of a` solution having the` composition indicatedin Table I was reduced bycontact with metallic ironfto a potential of+0.08 V. as measured with a platinum-standardcalomel electrode system.

24.7 grams per 'liter of calcium carbonate and 25 grams per liter ofaqueous hydrogen uoride added to the sointionV produced a .precipitateWeighing 694 grams,v sub statially free of vphosphate after washing,andcontaining 0.21% uranium and about 78% of the recoverable tluorine.'lhisuranium recovery amounts to abo'ut94`% bfglfon the content of theprecipitates and tiltrates,

. of 900 C.

esta, 1ers ing an excess of concentrated sulfuric acid and heating themixture for 130 minutes to a maximum temperature was leached with 325cc. of 0.2 N sulfuric acid at a temperature `of 90C. for 15 minutes witha resultant removal of 68% of the uranium bythe leach liquor. A

second Vleaching of the same material with 325 cc. of 0.4 N sulfuricacid for minutes at 95 C. raised the overall amount of uranium leachedby the two successive operations to 98% of the uranium originallypresent in the material being treated.

EXAMPLE `B 20.7 liters of this acid was reduced by circulation through ametallic iron packed column until the electromotive potential of thesolution as measured between platinum and standard calomel electrodeswas -|-0.033 V. The reduced solution was then treated with 360 g. ofcalcium carbonate and 476 g. of uorine as an aqueous solution of HF toprecipitate the uranium. After washing with water and drying at 120 C.for 24 hours the precipitate weighed 805 g. and had the compositionindicated in Table Ill.

This percentage of uranium amounts to `a recovery of 76% of the uraniumoriginally present in the acid and is slightly lower than that normallyobtained due to incomplete reduction of the acid. Portions of thisprecipitate were deiluorinated and calcined by heating with`concentrated sulfuric acid using the conditions indicated in Table lV.Portions of these deuorinated materials were then leached employing theconditions indicated under the appropriate heading of Table IV. Thewater and 0.092 N sulfuric acid leach being done on the material heatedfor minutes at 800 C. and the 0.18 N sulfuric acid leach being done onthe material heated for 17 0 minutes at 900 C.

Table IV DEFLUORINATION H1804 percent excess 6. 9 '5 Defluorinationtime, 100 170 Max. Temp. of dsuornatlon, "O 800 900 LEACH [50 g. ofdeuorinated material leached with 125 ce. of reagent.]

H1804 0. 092 N B leach grams of this Vdeuorinated material;

Uranium contained in these leach liquors was recovered by treating 25ml.portions with 0.5 g. of -sodium hydro sulte resulting in theprecipitation of 8l and 99.9%,V respectively, of the uranium fromleachliquors A and `C asa precipitate containing 13.4 and 19.1% of U.

' Also, the uranium was recovered as uranyl 4phosphate by adding alkalito the solutions. 25 ml. of leach solution B was treated with 26ml. of0.1 N NaOH resulting in the precipitation of 41%y of the uranium present'ein a precipitate containing 5.1% U. 25 ml.rv portiens'of solu tion Cwere treated with 18.8 and 39.2 ml. of 0.1 N NaOH, respectively,resulting in the precipitation of 99.7 and 99.9% of the uranium in aprecipitate containing 10.3 and 9.7% U.

EXAMPLE C 2 liters of 35% commercial phosphoric acid' containing 108Ing/liter of UsO was reduced with iron and the uranium precipitated bythe addition of 36 g. of calcium carbonate and 52.8 g. of .hydrogenfluoride. About 90% of the uranium was found in the precipitate. Thisprecipitate containing 190 mg. of U3O8 was successively leached withsodium carbonate solution as follows:

Extracted Solution:

A 30 g. NazCO; iu 500 m1. H30 at 25 C 3. 6 B,.-. 40 g. NazCOa in 500 m1.H20 211:80 C... 69.0 0...- 25 g. NazCOa in 500 m1. H00 at 80 C.-. 68. 4D 25 g. NazCOs in 500 ml. H20 at 80 0..-. 3G. 5

Total (mg. Usos) 177.5

120.7 g. of the solid remaining after leaching had the followingcomposition:

The four solutions obtained from above were then employed in successionto leach a second similarprecipitate weighing 68 g. and containing 136mg. U30?, by heating for 20 min. at 80 C. with the following results.

Usos content Usos Content. 'ofLeachng .After Leach- Solution, mg. Y ing,mg.

In the same manner another precipitatey was treated with these foursolutions resulting in the U3O8 content of the combined solutions beingraised to 500 mg. The combined solutions were heated with hydrochloricacid to remove the carbonate and ammonia was adde'd to precipitate theuranium. There was obtained a vprecipitate weighing 0.7338 g. whichYcontained 66.3% U3O8, .23.8% P04, 9.7% SOQ, and 0.55% vanadium, thusindicating that the product was fairly pure uranyl phosphate.

EXAMPLE D Both leached (with Na2CO3) and unleached precipitated materialwas deiiuorinat'ed by heating-with ammonium sulphate and sulfuric acida'sf indicated in Table V.

Table V E Fluorine (NHm'SOiy Time and Temp. Fluoride Wt. ofppt gContent, used, g. of heating sublim ed,

f .percent percent n.06 1 30. a 10. 02 24 lin/4 sa 2 5.00 49; 2 8. 201.5 hu] 97. 1 10.00 49. 2 16. 00 1.5 hr. 97. 9 10.00 44. 9 20. 00 1.5hr. 99. 1 15.00 44. 0 10. 01 1.5 hr. 08. 9 t1.98 44. 9 8. 01 v1.5 b'rx/98. 5 5.06 2 44. 9 10. 00 1.5 hr. 08. 8 5.04 2 44. 9 18. 00 1.5 lin/45090. 0 7.05 2 44. 9 3. 83 1.5,hr.l450 C-.-. 70. 8 5.01 2 44. 9 6. 02 1.5IIL/450 90. 1

4e 1 135 /i hr /s75jor 07.1y 46 7 135 %hr.'/ 0 0..-- 99:0 46 7 135 lv%h1./500G..'.; 94. 4

I 'Lcached Unleached. v

EXAMPLE E Variousreducing agents were employedto lreduce ltwoliterportions of `industrial (35%) phosphoric acid containing "1'08 mg.{130s/liter and the uranium was precipitated under the conditions'andwith the results 'indicated inl Table VI.

Table lVl UsOain Recovery Reducing Agent Preeltltatiug ppt. Wt., ppt.,perof UOs Agents g. cent fromacid,

g percent Y say' 0; 14" 72 s2 0. 2a s1 Uranium was recovered fromseveral of the precipitates listed in Table VI by the following methods:

Precipitate (9) weighing 82 g. and containing 0.22% or 180 mg. of U30Bwas heated with 400 m1. of afsolution containing 50 g. of Na2CO3 for 2hours with a resultant leaching of 128 mg. or 71.3% of the uranium.

Precipitate (10) a well washed precipitate weighing 187.8 g. wet whichis equivalent to 53.3 g. dry and'containing 0.27% or 144 mg. of U3O8 was'heated' for 11/2v hours with 400 ml. of a solution containing 50 g.NazCOa resulted inthe leaching of of'theuranium from;

the precipitate. This leach solution which now amounted to 595 ml. dueto wash water inclusion was lthen ernployed 4to leach the entire wetprecipitate l(l1)','after an additional 25 g. of Na2CO3 was added and'with heating for lil/ahours. The leach filtrate contained 259 mg. ofUBOB amounting to a 57% extractionv of uranium from the .secondprecipitate in addition to thatextract'ed from thefrs't.

Precipitate (12) fromabove was subjected tosuccessive` leachings with"fresh portions of carbonate solution as follows:

4(1)"1`l1fey entire Yprecipitater containing 190 mg'. of U'BCg was-stirred'for l1/2" hours with afcold'solution of 50 g. of Ninco, in 50omi. ofwater. The leaehura contained 3.6 mg. U30 K y (2) The residualprecipitate from (l) was stirred for 45 minutes with heating withr asolution of 40 g. of Na2CO3 in 500 ml. of water and theprecipitate wasseparated from the solution by filtering. This filtrate con-V tained 69mg. of USOS.

(3) The residual precipitate from 2, precedingwas heated with a 500 ml.of a solution containing 25 g. NagCO?,V for 1% hours andwas thenseparated from the solution by filtering. The iiltrate contained 68.4mg.

EXAMPLE F A variety of reducing agents were employed to reduce 100 ml.portions of 35% Ycommercial vphosphoric acid containing 10.6 mg. U3Oausingthe conditions and with the results ndicatedin Table VII. Thereduced uranium was then precipitated by the addition of 1 g. CaO

i and 2.64 g. of HF to the reduced solution while con-- tained in aplastic beaker. In some cases HF was addedA with the reducing agent witha` consequent improvement in the amount of uranium precipitated.Increased temperature also improves the recovery.

Table VII t'llme, Tempera- Recovery Reducing Agent hr. ture, C. UaOx,

percent 30 g. Pb l0 mesh 95 8l 10 g. Pb 10 mesl1 2 95 67 l g. Pb l0mesh.. 3 95 78 2 g. Bn 10 mesh..- )i 95 63 g. Sn 10 mes B l5 67 30 g. CuTurnings-.. 1 95 64 g. Cu Turniugs 1-- 1 95 88 2 g. Cu 80 mesh shot. 195 76 2 g. Cu 8O mesh shot 95 86 2 g. Fe Turnings...-- 2 25 94 2 g. FeTurnings 1 95 83 5 g. Fe Turnings-. 1 25 S9 4 g. Zu 20 mesh--.. 1 25 812 g. Zn 20 mesh 1 1 25 94 1 g. Zn Amalgarn 1 25 75 10 g. Zn Amalgam- 425 89 5 g. Zn Amalgam.. 1 25 8S 2 g. Zn Amalgam 1 25 88 1 2.47 g. HFadded.

In the experiments using copper, only small amounts of copper appearedto dissolve. Zinc and zinc amalgam were effective reducing agents withonly small amounts although zinc dissolved with evolution of hydrogen.Hydrogen is not evolved when lead is used.

In the foregoing, the uranium content is indicated in amounts orpercentages of USOS. However, it is not intended that this terminologyis to be construed literally but is intended to indicate that theequivalent quantity of uranium is present in the appropriate form asindicated by context. v

While there has been described in the foregoing what may be consideredpreferred embodiments of the invention, itis believed that variousmodifications may be made therein without departing from the spirit andscope of the invention and it is intended to cover all such as fallwithin the scope of the appended claims.

What is claimed is:

1. The method of recovering uranium values from a phosphatic solutioncontaining impurity materials the iluorides of which are insoluble insaid phosphatic solution comprising reducing the uranium to the uranousstate in the solution, adding a soluble calcium coprecipitant compoundto said solution, adding a soluble uoride to said solution, whereby saidcoprecipitant and some of said impurities are coprecipitated withtheuranium to form complex' uoride precipitate, leaching uranium fromthe complex tluoride precipitate, and recovering uranium from the leachsolution.

2. The method of recoving uranium values from a crude phosphaticsolution containing materials the fluoridesA of which are insoluble insaid phosphatic solution, comprising reducing the uranium to the uranousstate in the solution, adding -a soluble calcium coprecipitant compoundto said solution, adding a soluble uoride to said solution, whereby saidcoprecipitant and the uranium are coprecipitated to form a complexuoride precipitate, de-

liuorinating and calcining Vsaid precipitate, leaching uranium from saiddelluorinated and calcined precipitate, and recovering uranium from theleach solution.

3. The method of recovering uranium values from a crude phosphaticsolution comprising reducing the uranium to theluranous state in thesolution, adding a soluble calcium coprecipitant compound to saidsolution, adding a soluble uoride to said 'solution whereby the uraniumis coprecipitated'with said coprecipitant to form a com plex lluorideprecipitate, leaching uranium from said complex fluoride precipitateunder acidic conditions, and

`recovering uranium from the leach solution.

precipitating said uranous uranium from the solution by adding solubleuorides and at least one coprecipitant selected from the groupconsisting of calcium, barium, magnesium, aluminum, sodium, potassiumand ammonia which lform insoluble lfluoride insaid solution thereto,leaching uranium from said precipitate With an acid solution, wherebysome phosphate is also leached, and treating said acidic solution with areducing agent to precipitate said uranium from the leach solution asuranous phosphate.

5. The process for recovering uranium from acidic. phosphatic solutionscomprising reducing said uranium to the uranous state in the solution,precipitating said uranous uranium from the solution by adding theretosoluble uorides and coprecipitants selected from the group consisting ofcalcium, barium, magnesium, aluminum, sodium, potassium and ammoniawhich form in soluble fluorides in said solution, deuorinating andcalcining said precipitate by heating with a material selected from thegroup consisting of sulfuric acid and ammonium sulfate, leaching uraniumfrom said deuorinated and calcined precipitate with an acidic aqueoussolution, and recovering uranium from said acidic leach solution.

6. The method of recovering uranium from an acidic phosphatic solutioncomprising reducing said uranium to the uranous state in solution,precipitating said uranous uranium by adding to the solution solublelluorides and coprecipitants -selected from the group consisting ofcalcium, barium, magnesium, aluminum, sodium, potassium and ammoniawhich form insoluble tluorides in said solution, leaching uranium fromsaid precipitate with a carbonate solution, acidifying said leachsolution, and precipitating the uranium from the acidited leach solutionas uranium diuranate by the addition of ammonia.

7. The meth-od ofI recovering uranium from a phosphatic solutioncomprising reducing said uranium to the uranous state in the solution,precipitating said uranous uranium by adding to said solution a solubleuoride and coprecipitants selected from the group consisting of calcium,barium, magnesium, aluminum, sodium, potassium and ammonia whchforminsoluble lluorides in said solution, leaching uranium from saidprecipitate with an acid solution, and precipitating the uranium fromsaid solution as uranyl phosphate by adjusting the pH of the solutionwith an alkaline material.

8. The process as defined in claim 7, wherein said acid is a materialselected from the group consisting of hydrochloric, nitric, citric,tartaric, acetic and boric acids.

9. The process of recovering uranium from an acidic phosphatic Solutioncontaining. impurities. including phosphate, sulfate, iron,l silicon,vanadiumand titaniunncomprising reducing the uranium to the uranousstate in thev solution, adding a soluble calcium coprecipitant compound`to the solution,lt reatiug said solution with a soluble uoride toprecipitate said; material and at least some of said impurities from thesolutionwhereby the uranium is coprecipitated with the. precipitate,leaching uranium from said precipitate with a material selectedfrom thefgroup consisting of sodiunieearbonate and ammonium carbonate, acidifyingthe leach solution, and precipitating the uranium from the acidiiiedsolution withy ammonia.

l0. The process fortrecovering uranium from anacdic phosphatic solutioncontaining impuritiesfincluding phosphate, sulfate, iron,` silicon,vanadium, and titanium, com prising reducing the uranium to the-uranousstate in the solution, adding a soluble calcium coprecipitant compoundto the solution, treatingI Said solution with a soluble uoride toprecipitate said material and at least some of said impurities fromthesolution, whereby the uranium is coprecipitated4 with theprecipitate, detluorinating and calcining said precipitateby heatingwith a material selectedfrom theA group consisting of. sulfuric acid andammonium sulfate, leaching uranium from the deuorinated and caleinedprecipitate with amaterial selected from the group consisting of water,sulfuric acid and nitric acid, whereby some phosphate is also leached,andn precipitating uranyl` phosphate by making. the solution alkaline.

1l. The processfor recovering uranium from an acidic phosphatic solutioncontaining impurities 'including phosphate, sulfate, iron, silicon,vanadium, and titanium, com.- prising reducing the uranium to theuranous statein the solution, adding aI soluble calcium coprecipitantcom-l pound to the solution, treating saidsolutionwith a solubleuorideto precipitate.said'material andat least some of said impuritiesfrom the solution, whereby the uranium isl eopnecipitatedwiththeflprecipitate, defluorinating and calcining said precipitate byheatiugvvwith a material selectedfrom the; group-consisting of sulfuricacid and ammonium sulfate, leaching, uranium froml thedeuorinated andcalcined precipitate with a material selected from theA groupconsisting, of water, sulfuric acid and nitric acid, whereby somephosphate is also leached, and precipitating` uranous4 phosphate fromsaid solution by reducingthe uranium tothe uranousstate.

12. The process for recoveringuranium from an acidic phosphatic-vsolution comprising; reducing the uranium to the uranous state insolution, adding to said solution copriecipitaut. materials selectedvfrom the group consisting of -calcium,vbarium, I naguesium, aluminum,sodium, potassiurn and ammoniaV which form. insoluble fluorides in saidsolution,treatingsaidl solution with soluble uorides to, precipitatesaid coprecipitantmaterials, whereby the uranium is coprecipitatedthenewithJcachingthe uranium from the precipitate with an, acidicsolution, whereby some phosphate is also leached, and precipitatinguranous phosphate from the, solution; by. reducing the uranium to theuranous state.

1.3. The process as defined in claim l2, wherein said acidic solutionwhich is, employed toleach't'he precipitate is amaterialselected-fromme.- group consisting of hydrochloric, nitric;` citric,tartaric, acetic andboric acid solutions; t

1,4. ThevProcoSs as.A deinedf inf claim l0 wherein said soluble.lluQride is ai material; selected, from the group consisting of sodium,potassium, barium and, ammonia fluorides.

References Citediny the le of this patent Mellor: InorganicyandTheoretical.Chemistry, vol. 12, pageA 74 (1932). Publ. by Longmans,Green & Co., London;

1. THE METHOD OF RECOVERING URANIUM VALUE FROM A PHOSPHATIC SOLUTIONCONTAINING IMPURITY MATERIALS THE FLUORIDES OF WHICH ARE INSOLUBLE INSAID PHOSPHATIC SOLUTION COMPRISING REDUCING THE URANIUM TO THE URANOUSSTATE IN THE SOLUTION, ADDING A SOLUBL E CALCIUM COPRECIPITANT COMPOUNDTO SAID SOLUTION, ADDING A SOLUBLE FLUORIDE TO SAID SOLUTION,WHERBY SAIDCOPRECIPITATED AND SOME OF SAID IMPURITIES ARE COPRECIPITATED WITH THEURANIUM TO FORM A COMPLEX FLUORIDE PRECIPITATE, LEACHING URANIUM FROMTHE